JP2002285253A - Method for leaching zinc concentrate - Google Patents
Method for leaching zinc concentrateInfo
- Publication number
- JP2002285253A JP2002285253A JP2001091738A JP2001091738A JP2002285253A JP 2002285253 A JP2002285253 A JP 2002285253A JP 2001091738 A JP2001091738 A JP 2001091738A JP 2001091738 A JP2001091738 A JP 2001091738A JP 2002285253 A JP2002285253 A JP 2002285253A
- Authority
- JP
- Japan
- Prior art keywords
- leaching
- zinc
- zinc concentrate
- iron ions
- grinding
- Prior art date
- Legal status (The legal status is an assumption and is not a legal conclusion. Google has not performed a legal analysis and makes no representation as to the accuracy of the status listed.)
- Granted
Links
Classifications
-
- Y—GENERAL TAGGING OF NEW TECHNOLOGICAL DEVELOPMENTS; GENERAL TAGGING OF CROSS-SECTIONAL TECHNOLOGIES SPANNING OVER SEVERAL SECTIONS OF THE IPC; TECHNICAL SUBJECTS COVERED BY FORMER USPC CROSS-REFERENCE ART COLLECTIONS [XRACs] AND DIGESTS
- Y02—TECHNOLOGIES OR APPLICATIONS FOR MITIGATION OR ADAPTATION AGAINST CLIMATE CHANGE
- Y02P—CLIMATE CHANGE MITIGATION TECHNOLOGIES IN THE PRODUCTION OR PROCESSING OF GOODS
- Y02P10/00—Technologies related to metal processing
- Y02P10/20—Recycling
Landscapes
- Manufacture And Refinement Of Metals (AREA)
Abstract
Description
【０００１】[0001]
【発明の属する技術分野】本発明は、亜鉛、更には鉛、
金、銀等の有価金属元素を含有する硫化物である亜鉛精
鉱（単に、亜鉛精鉱という。）から亜鉛、更には鉛、
金、銀等の有価金属元素および単体硫黄を回収する湿式
亜鉛製錬における亜鉛精鉱の浸出法に関する。[0001] The present invention relates to zinc, and further to lead,
From zinc concentrate (simply called zinc concentrate), which is a sulfide containing valuable metal elements such as gold and silver, zinc, and further lead,
The present invention relates to a method for leaching zinc concentrate in a wet zinc smelting for recovering valuable metal elements such as gold and silver and elemental sulfur.
【０００２】[0002]
【従来の技術】亜鉛精鉱の浸出等に関する従来の技術と
しては、特許公報第２８５６９３３号および特公平６−
４３６１９に開示された方法がある。まず、特許公報第
２８５６９３３号によれば、次に述べるような方法を用
いて亜鉛精鉱が処理される。すなわち、亜鉛精鉱の浸出
は二段階で行われるものであって、亜鉛精鉱を焙焼して
亜鉛▲か▼焼物を生成後、中性浸出を行う。次いで電解
処理工程において得られた戻し酸を用いて強酸浸出を行
い、未浸出亜鉛精鉱及び焙焼により生成した難溶性のジ
ンクフェライトを分解する。亜鉛の浸出に必要な三価の
鉄イオンはジンクフェライト分解によって生じる鉄量だ
けでは不十分のため、浸出後液中の二価の鉄イオンを酸
化して再利用するという方法により浸出を行っている。
この結果９０〜９５℃において６〜１０時間かけて９９
％の亜鉛回収率を達成している。また、浸出時に生成す
る残渣は、溶鉱炉を用いて乾式冶金処理して有価金属を
回収するか、若しくは浸出残渣を浮選にかけて有価金属
を濃縮し回収している。2. Description of the Related Art Japanese Patent Publication No. 2856933 and Japanese Patent Publication No.
There is a method disclosed in US Pat. First, according to Patent Publication No. 2856933, zinc concentrate is treated using the method described below. In other words, the leaching of zinc concentrate is performed in two stages, and after the zinc concentrate is roasted to produce a zinc- or calcined product, neutral leaching is performed. Next, strong acid leaching is performed using the returned acid obtained in the electrolytic treatment step to decompose the unleached zinc concentrate and the hardly soluble zinc ferrite generated by roasting. Since the trivalent iron ions required for zinc leaching are not enough by the amount of iron generated by decomposition of zinc ferrite, leaching is performed by oxidizing and reusing divalent iron ions in the solution after leaching. I have.
As a result, 99 to 99 hours at 90 to 95 ° C.
% Zinc recovery. In addition, the residue generated during leaching is subjected to dry metallurgy in a blast furnace to recover valuable metals, or the leaching residue is subjected to flotation to concentrate and recover valuable metals.
【０００３】一方、特公平６−４３６１９に記載されて
いる亜鉛精鉱の処理方法は、亜鉛精鉱を少なくとも２段
階以上にわたって浸出する方法であり、鉱石を粉砕して
微粒化した後、第１段階浸出では、温度１２５〜１６０
℃、最終遊離硫酸濃度２０〜６０ｇ／L、三価の鉄イオ
ン濃度１〜５ｇ／Lとなるように酸素圧をかけた状態で
加圧浸出を行い、亜鉛を不完全溶解する。その後の第２
段階浸出では、大気圧下において、電解処理工程で発生
する戻り酸を過剰に用い、遊離硫酸濃度６０〜１６０ｇ
／L、三価の鉄イオン濃度２〜３ｇ／Lとなるように酸素
を供給した状態で浸出を行い、硫酸亜鉛溶液と浸出残渣
を形成させる。この際形成される残渣には残留亜鉛、
銅、鉄、大部分の鉛及び貴金属が含まれているので、浮
選により分離回収を実施する。On the other hand, a method for treating zinc concentrate described in Japanese Patent Publication No. 6-43619 is a method of leaching zinc concentrate in at least two or more stages. In stage leaching, temperatures between 125 and 160
C., pressurized leaching is performed under oxygen pressure so that the final free sulfuric acid concentration is 20 to 60 g / L and the trivalent iron ion concentration is 1 to 5 g / L, and zinc is incompletely dissolved. Then the second
In the step leaching, under atmospheric pressure, excess return acid generated in the electrolytic treatment step is used, and the free sulfuric acid concentration is 60 to 160 g.
/ L, leaching is performed in a state where oxygen is supplied so as to have a trivalent iron ion concentration of 2 to 3 g / L to form a zinc sulfate solution and a leaching residue. The residue formed at this time is residual zinc,
Since it contains copper, iron, most of lead and precious metals, separation and recovery are performed by flotation.
【０００４】[0004]
【発明が解決しようとする課題】上に述べた従来の方法
は、既存の焙焼−浸出−電解工程への組み込みが可能で
あり、かつ既存の設備への増強が少なくて良いという利
点があった。また、投入する亜鉛精鉱中の亜鉛の回収率
も高く、かつ銅、鉛、および他の貴金属の回収も同時に
行うことが可能であるという優れた点がある。しかし、
亜鉛を溶液中へ完全に溶解するのに要する時間が長いこ
とや、酸化反応を促進させるためにオートクレーブ等の
圧力容器や規模の大きい反応槽等を必要とするため、建
設費が高いという問題を抱えていた。The above-mentioned conventional method has the advantages that it can be incorporated into the existing roasting-leaching-electrolysis process, and that there is little need to enhance existing equipment. Was. Further, there is an excellent point that the recovery rate of zinc in the zinc concentrate to be fed is high, and it is possible to simultaneously recover copper, lead and other noble metals. But,
It takes a long time to completely dissolve zinc in the solution, and requires a pressure vessel such as an autoclave or a large-scale reaction tank to promote the oxidation reaction, which results in high construction costs. I was holding it.
【０００５】[0005]
【課題を解決するための手段】このような従来技術の問
題を解決するために本発明者らは種々の検討を重ねた結
果、亜鉛精鉱を磨鉱しながら浸出することにより、又は
磨鉱と浸出を別個に行うことによっても、該亜鉛精鉱の
表面にある浸出反応を阻害する成分を剥離または分離除
去することで、亜鉛の浸出速度が飛躍的に向上すること
を発見した。また、亜鉛精鉱を浸出する際にポンプで液
を循環させている間に酸素を配管内に供給し、配管内を
加圧状態にすることで、反応により消費されて二価の鉄
イオンに変化した鉄イオンを三価の鉄イオンに再生して
浸出に再利用することができ、大規模な圧力容器を用い
ずに鉄イオンを酸化させることに成功した。The inventors of the present invention have made various studies in order to solve the problems of the prior art. As a result, the zinc concentrate is leached while grinding, or It has also been found that the leaching rate of zinc is dramatically improved by separating or removing components that inhibit the leaching reaction on the surface of the zinc concentrate by separately performing leaching with the zinc concentrate. In addition, oxygen is supplied into the pipe while circulating the liquid with a pump when leaching zinc concentrate, and the pipe is pressurized, so that it is consumed by the reaction and converted into divalent iron ions. The changed iron ions can be regenerated into trivalent iron ions and reused for leaching, and the iron ions can be oxidized without using a large-scale pressure vessel.
【０００６】すなわち本発明は、第１に、遊離硫酸と三
価の鉄イオンとを含有する水溶液中において亜鉛精鉱を
磨鉱することによって該亜鉛精鉱中の亜鉛を浸出するこ
とを特徴とする亜鉛精鉱の浸出法；第２に、前記磨鉱を
大気圧下で行う、第１記載の浸出法；第３に、前記浸出
に伴って前記亜鉛精鉱の粒子表面に生成される副生物を
前記磨鉱により剥離して該浸出時間を短縮する、第１ま
たは２記載の浸出法；第４に、前記浸出により前記三価
の鉄イオンが還元されて生じた二価の鉄イオンを含有す
る浸出後液中に酸素を供給することによって該二価の鉄
イオンを三価の鉄イオンに酸化した液を前記磨鉱の工程
に繰り返す、第１〜３のいずれかに記載の浸出法；第５
に、前記磨鉱の工程へ供給または繰り返される前記水溶
液または前記浸出後液の送液配管内に酸素を供給して該
管内を加圧状態にする、第１〜４のいずれかに記載の浸
出法；第６に、前記水溶液中の遊離硫酸濃度を浸出終了
時点で４０ｇ／L以上としてジャロサイトの生成を抑制
する、第１〜５のいずれかに記載の浸出法；第７に、前
記水溶液中の三価の鉄イオン濃度を５〜６０ｇ／Lの範
囲とする、第１〜６のいずれかに記載の浸出法；第８
に、鉛、金および銀のうちの少なくとも１種の金属元素
を含有する前記亜鉛精鉱中の該金属元素が濃縮された前
記浸出後の残渣を浮選し該金属元素を回収する、第１〜
７のいずれかに記載の浸出法；第９に、前記浮選によっ
て得られた浮鉱を硫黄の融点以上の温度に加熱して該浮
鉱中の単体硫黄を揮発回収する、第８記載の浸出法、を
提供する。That is, the present invention is characterized in that firstly, zinc in a zinc concentrate is leached by grinding the zinc concentrate in an aqueous solution containing free sulfuric acid and trivalent iron ions. Leaching method for zinc concentrate; secondly, performing the grinding operation at atmospheric pressure; leaching method according to the first aspect; thirdly, secondary leaching generated on the surface of particles of the zinc concentrate along with the leaching. The leaching method according to claim 1 or 2, wherein the leaching time is shortened by exfoliating the organisms with the grinding, and fourthly, divalent iron ions produced by reducing the trivalent iron ions by the leaching are produced. The leaching method according to any one of claims 1 to 3, wherein a liquid obtained by oxidizing the divalent iron ions into trivalent iron ions by supplying oxygen to the leaching liquid is repeated in the grinding step. The fifth
The leaching according to any one of claims 1 to 4, wherein oxygen is supplied to a pipe for feeding the aqueous solution or the liquid after leaching, which is supplied or repeated to the grinding process, so that the pipe is pressurized. Sixth, the leaching method according to any one of the first to fifth aspects, wherein the concentration of free sulfuric acid in the aqueous solution is set to 40 g / L or more at the end of leaching to suppress jarosite formation; The leaching method according to any one of the first to sixth aspects, wherein the trivalent iron ion concentration in the medium is in the range of 5 to 60 g / L;
A flotation of the residue after leaching in which the metal element in the zinc concentrate containing at least one metal element of lead, gold and silver is concentrated to recover the metal element; ~
9. The leaching method according to any one of 7) to 9), wherein the flotation obtained by the flotation is heated to a temperature equal to or higher than the melting point of sulfur to volatilize and recover elemental sulfur in the flotation. Leaching method.
【０００７】[0007]
【発明の実施の形態】亜鉛精鉱を電解処理により発生し
た遊離硫酸濃度１５０〜２００ｇ／L程度の電解工程の
戻り酸および鉄を除去した後に発生する后液を用いて８
０〜９５℃まで昇温させ亜鉛の浸出反応を起こす。この
際の反応は以下の通りである。 ZnS + Fe2(SO4)3 → ZnSO4 + 2FeSO4 + S ・・・Ａ式 Ａ式の反応を促進するために必要な三価の鉄は、処理す
る亜鉛精鉱中の鉄を用いることとする。その場合、浸出
時の三価の鉄イオン濃度は好ましくは５〜６０ｇ／Ｌの
範囲、さらに好ましくは５〜１５ｇ／Lとする。三価の
鉄イオン濃度が５ｇ／Ｌ未満ではＡ式の反応速度が不十
分であり、６０ｇ／Ｌ以上では効果が飽和してくる。こ
の反応は、時間と共に進行するものの、反応により生成
した単体硫黄等の副生物が亜鉛精鉱の粒子表面に生成、
付着するために反応界面が減少し、その結果その後の反
応速度が低下し、全ての亜鉛を浸出させるためには多大
の時間を要していた。本発明の方法では、Ａ式の反応で
生成し、粒子表面に生成、付着した単体硫黄等の副生物
を剥離または分離させるためにボールミル等の粉砕機を
用いて、亜鉛精鉱を磨鉱しながら浸出を行った。また、
磨鉱工程と浸出工程を別個に行うこともできる。すなわ
ち、亜鉛精鉱を一旦浸出した後に、磨鉱して粒子表面に
生成、付着した副生物を剥離、分離してから、再度浸出
し、引き続き磨鉱を行うというように交互に各工程を繰
り返すことによっても目的を達成することができる。な
お、磨鉱に使用する粉砕機は、亜鉛精鉱の粒子表面の単
体硫黄等副生物を剥離または分離する目的に適する装置
であれば、ボールミルに限定されない。このような粉砕
機には、例えばロッドミル、タワーミル、振動ミル、ア
トリションミル等がある。また、磨鉱により亜鉛精鉱が
より微細化され比表面積が増えるためより浸出を促して
いる。磨鉱前の亜鉛精鉱の粒度は、特に問わないが、よ
り浸出時間を短縮するためには小さい方が望ましく、好
ましくはメジアン径が１〜１００μｍ、９０％粒子径が
５０〜１０００μｍである。メジアン径が１μｍより細
かいと鉱石の移送時に飛散しやすくなり原料歩留まりの
低下を招く、１００μｍより大きいと効果が十分に得ら
れない。また、磨鉱条件としては磨鉱時のスラリー濃度
は高い方が良く、好ましくは３０ｇ/L以上が良い。BEST MODE FOR CARRYING OUT THE INVENTION Zinc concentrate is prepared by using a back solution generated after removing a return acid and iron in an electrolysis step of a free sulfuric acid concentration of about 150 to 200 g / L generated by electrolytic treatment.
The temperature is raised to 0 to 95 ° C. to cause a zinc leaching reaction. The reaction at this time is as follows. ZnS + Fe 2 (SO 4 ) 3 → ZnSO 4 + 2FeSO 4 + S ・ ・ ・ Formula A The trivalent iron required to promote the reaction of formula A is the iron in the zinc concentrate to be treated. And In this case, the concentration of trivalent iron ions at the time of leaching is preferably in the range of 5 to 60 g / L, and more preferably 5 to 15 g / L. When the trivalent iron ion concentration is less than 5 g / L, the reaction rate of the formula A is insufficient, and when the concentration is 60 g / L or more, the effect is saturated. Although this reaction proceeds with time, by-products such as elemental sulfur generated by the reaction are generated on the surface of the zinc concentrate particles,
The deposition reduced the reaction interface and consequently slowed the rate of the reaction, requiring a great deal of time to leach out all the zinc. In the method of the present invention, zinc concentrate is crushed using a pulverizer such as a ball mill to separate or separate by-products such as elemental sulfur generated and adhered to the particle surface, which are generated by the reaction of Formula A. While leaching. Also,
The grinding and leaching steps may be performed separately. That is, after the zinc concentrate is leached once, each step is alternately repeated, such as grinding and stripping off and separating by-products generated and adhered to the particle surface, leaching again, and subsequently grinding. This can also achieve the purpose. The pulverizer used for grinding is not limited to a ball mill as long as it is a device suitable for separating or separating by-products such as elemental sulfur on the surface of zinc concentrate. Such a pulverizer includes, for example, a rod mill, a tower mill, a vibration mill, an attrition mill, and the like. Further, the zinc concentrate is refined by the grinding and the specific surface area is increased, so that leaching is promoted. The particle size of the zinc concentrate before grinding is not particularly limited, but is preferably small in order to further shorten the leaching time, and preferably has a median diameter of 1 to 100 µm and a 90% particle diameter of 50 to 1000 µm. When the median diameter is smaller than 1 μm, the ore is liable to be scattered at the time of transfer, thereby lowering the raw material yield. When it is larger than 100 μm, the effect cannot be sufficiently obtained. As for the grinding conditions, the higher the slurry concentration at the time of grinding, the better, preferably 30 g / L or more.
【０００８】この浸出反応により浸出残渣が発生する
が、浸出条件によっては、反応時に鉛ジャロサイトが生
成する。この鉛ジャロサイトが存在すると、生成する浸
出残渣量が増大するために、残渣処理にかかるコストの
増大に繋がる。従って、浸出反応時にはジャロサイト生
成を抑制するために浸出終了時点での遊離硫酸濃度を４
０ｇ／L以上にする必要がある。次にＡ式の反応を見れ
ば明らかなように、亜鉛精鉱の浸出が進行するに伴い、
浸出に必要な三価の鉄イオンが消費され減少してくる。
三価の鉄イオンがなくなればＡ式の反応は進行せず、浸
出反応が停止する。これを防ぐための方法として、亜鉛
精鉱中の亜鉛量に相当する量の三価の鉄イオンを繰り返
すか、反応により発生した二価の鉄イオンを酸化するこ
とで三価の鉄イオンを再生させ、再利用する方法があ
る。この酸化反応は次のＢ式に示す通りである。 2FeSO4 + 1/2O2 + H2SO4 → Fe２(SO4 )３＋H2O・・・Ｂ式 鉄イオンの酸化反応を大気圧下で行うと、この酸化反応
は非常に速度が遅い。そこで、オートクレーブなどの圧
力容器を使用して加圧状態とし、反応速度を速めて酸化
するのが一般的である。しかし、オートクレーブ等の圧
力容器は扱いにくく、かつ高価な設備であるために、本
発明では加圧状態を配管内で作り出し、配管内を液が流
れる間に二価の鉄イオンの酸化を行うことで三価の鉄イ
オンを再生し、浸出工程へ戻して再利用する。[0008] Leaching residue is generated by the leaching reaction, but lead jarosite is generated during the reaction depending on leaching conditions. The presence of the lead jarosite leads to an increase in the amount of leach residue generated, which leads to an increase in cost for residue treatment. Therefore, the concentration of free sulfuric acid at the end of leaching should be 4
It is necessary to be 0 g / L or more. Next, as is clear from the reaction of the formula A, as the leaching of the zinc concentrate progresses,
Trivalent iron ions required for leaching are consumed and decrease.
When the trivalent iron ions disappear, the reaction of Formula A does not proceed, and the leaching reaction stops. To prevent this, trivalent iron ions are regenerated by repeating trivalent iron ions in an amount corresponding to the amount of zinc in the zinc concentrate, or by oxidizing divalent iron ions generated by the reaction. There is a way to make it reuse. This oxidation reaction is as shown in the following formula B. 2FeSO 4 + 1 / 2O 2 + H 2 SO 4 → Fe 2 (SO 4 ) 3 + H 2 O ・ ・ ・ Formula B When the oxidation reaction of iron ions is carried out under atmospheric pressure, this oxidation reaction is very slow. . Therefore, it is general to use a pressure vessel such as an autoclave to pressurize the reaction vessel, thereby increasing the reaction rate and oxidizing the reaction vessel. However, since pressure vessels such as autoclaves are difficult to handle and expensive equipment, in the present invention, a pressurized state is created in the pipe, and oxidation of divalent iron ions is performed while the liquid flows in the pipe. To regenerate the trivalent iron ions and return to the leaching process for reuse.
【０００９】以上のような浸出、鉄イオンの酸化を段階
的に実施することにより約３０分で亜鉛精鉱中の亜鉛分
の約９５％を浸出させることが可能となり、従来の方法
での反応時間を大幅に短縮することが可能となった。ま
た、浸出用の水溶液中において磨鉱することによってこ
の効果が得られるが、前述のように磨鉱と浸出とを交互
に行うことによってもこの効果を得ることができる。浸
出により得られた浸出後の液は浄液工程を経て、電解処
理工程へと送液されて液中から亜鉛が電気亜鉛として回
収される。また、浸出残渣中には鉛、銀、単体硫黄及び
その他の貴金属が混入しているためにこれらを分離除去
する必要がある。そのため、浸出残渣を浮選工程へと送
り、硫化物及び単体硫黄とその他の金属を分離し処理す
る必要がある。この場合、浸出後に固液分離操作を行っ
て浸出残渣を濃縮スラリーとした後、空気を吹き込みな
がら実液のまま浮選を行う。これにより、硫黄及び硫化
物は浮鉱側へ、鉛、珪酸および貴金属は尾鉱側へと移行
する。得られた浮鉱中の単体硫黄は融点以上の温度で揮
発させて気体とさせて、冷却回収する。尾鉱には鉛およ
び貴金属が含まれているので、乾式冶金処理によりこれ
らの鉛、貴金属を回収する。By leaching and oxidizing iron ions in a stepwise manner as described above, about 95% of the zinc content in the zinc concentrate can be leached in about 30 minutes. Time has been greatly reduced. This effect can be obtained by grinding in an aqueous solution for leaching, but this effect can also be obtained by alternately performing grinding and leaching as described above. The leached liquid obtained by leaching is sent to an electrolytic treatment step through a liquid purification step, and zinc is recovered from the liquid as electric zinc. Further, since lead, silver, elemental sulfur and other noble metals are mixed in the leaching residue, it is necessary to separate and remove these. Therefore, it is necessary to send the leaching residue to a flotation step to separate and process sulfide and elemental sulfur from other metals. In this case, after performing the solid-liquid separation operation after leaching, the leaching residue is made into a concentrated slurry, and then flotation is performed with the actual liquid while blowing air. Thereby, sulfur and sulfide move to the flotation side, and lead, silicic acid and precious metal move to the tailing side. The obtained elemental sulfur in the flotation is volatilized at a temperature equal to or higher than the melting point to form a gas, which is cooled and recovered. Since tailings contain lead and precious metals, these lead and precious metals are recovered by pyrometallurgical treatment.
【００１０】[0010]
【実施例】［実施例1］ 粉砕機として、市販の試験用
アトリションミル型粉砕機アトライター（商品名）を使
用した。アトライターの容量は５．４L（２００ｍｍφ
×１７６ｍｍＨ）、材質はＳＵＳ３０４、モーター回転
数は１７０ｒｐｍ以上、使用ボールはアルミナボール
（９ｍｍφ、約３ｋｇ投入）である。浸出用水溶液とし
て、亜鉛濃度を１００ｇ／L、三価の鉄イオン濃度を３
０ｇ／L、遊離硫酸濃度を４０ｇ／Lに調整した水溶液を
用意した。亜鉛精鉱は、表１の組成を有する亜鉛精鉱を
使用した。Ｚｎ、Ｆｅ、Ｐｂ、Ｃｄ、Ｃｕ等の金属元素
は、硫化物の形で亜鉛精鉱中に含有されている。亜鉛精
鉱の粒度は、メジアン径が２５μｍ、９０％粒子径が７
０μｍである。EXAMPLES Example 1 A commercially available attrition mill type pulverizer attritor (trade name) was used as a pulverizer. The capacity of the attritor is 5.4L (200mmφ
× 176 mmH), the material is SUS304, the motor rotation speed is 170 rpm or more, and the balls used are alumina balls (9 mmφ, about 3 kg charged). As an aqueous solution for leaching, a zinc concentration of 100 g / L and a trivalent iron ion concentration of 3
An aqueous solution adjusted to 0 g / L and the concentration of free sulfuric acid to 40 g / L was prepared. As the zinc concentrate, a zinc concentrate having a composition shown in Table 1 was used. Metal elements such as Zn, Fe, Pb, Cd, and Cu are contained in zinc concentrate in the form of sulfide. The particle size of the zinc concentrate has a median diameter of 25 μm and a 90% particle diameter of 7%.
0 μm.
【００１１】[0011]
【表１】 [Table 1]
【００１２】上記の浸出用水溶液２．０Ｌを上記の粉砕
機の中に入れ、９０℃まで昇温した。昇温した上記水溶
液に上記の亜鉛精鉱６０ｇを添加して、粉砕機を運転さ
せて磨鉱、浸出を開始した。５分毎に粉砕機中のスラリ
ーをサンプリングしながら３０分間粉砕機の運転を続
け、粉砕機内で亜鉛精鉱と上記水溶液とを反応させた。
採取した各サンプルを濾過し、ケーキ（残渣）を十分水
洗した後、残渣品位を測定し、亜鉛精鉱品位と残渣品位
から亜鉛浸出率を求めた。上記の条件で試験を実施した
結果、表２に示すような亜鉛浸出率が得られ、僅か３０
分の浸出で９５％の亜鉛浸出率を達成できることが確認
された。[0012] 2.0 L of the above aqueous solution for leaching was put into the above-mentioned pulverizer and heated to 90 ° C. 60 g of the above zinc concentrate was added to the heated aqueous solution, and a grinder was operated to start grinding and leaching. The operation of the crusher was continued for 30 minutes while sampling the slurry in the crusher every 5 minutes, and the zinc concentrate and the aqueous solution were reacted in the crusher.
Each collected sample was filtered, and the cake (residue) was thoroughly washed with water. The residue quality was measured, and the zinc leaching rate was determined from the zinc concentrate quality and the residue quality. As a result of performing the test under the above conditions, the zinc leaching rate as shown in Table 2 was obtained, and only 30%
It was confirmed that 95% zinc leaching rate could be achieved by minute leaching.
【００１３】[0013]
【表２】 [Table 2]
【００１４】[実施例２] 実施例１に示したアトライタ
ーを用いて磨鉱、浸出試験を行った。試験は亜鉛濃度
１００ｇ/L、三価の鉄イオン濃度 １５g/L、遊離硫酸濃
度 ４０g/Lの母液を作成し、この母液の中に亜鉛精鉱を
前記母液に対して３０g/Lとなるように添加し、１０分
間浸出を行った。浸出後、スラリーを２００ g/Lに調整
し、アトライターで１分間粉砕を行った。粉砕後、スラ
リーを濾過し、残渣を得た。ここまでの手順を１回の試
験とし、その後残渣を母液に入れ、再度繰り返し試験を
行った。この繰り返し操作回数と亜鉛浸出率の関係を調
査した。この際の亜鉛精鉱は実施例１と同一組成のもの
を用いた。その結果を以下の表３に示す。[Example 2] Using the attritor shown in Example 1, a grinding and leaching test was performed. Test zinc concentration
A mother liquor of 100 g / L, trivalent iron ion concentration of 15 g / L and free sulfuric acid concentration of 40 g / L was prepared, and zinc concentrate was added to the mother liquor so as to be 30 g / L with respect to the mother liquor. Leaching was performed for 10 minutes. After leaching, the slurry was adjusted to 200 g / L, and pulverized for 1 minute with an attritor. After grinding, the slurry was filtered to obtain a residue. The procedure up to this point was defined as one test, and the residue was then placed in mother liquor, and the test was repeated again. The relationship between the number of repetitive operations and the zinc leaching rate was investigated. The zinc concentrate used herein had the same composition as in Example 1. The results are shown in Table 3 below.
【００１５】[0015]
【表３】 [Table 3]
【００１６】このように上述の操作を２回繰り返し、最
後に１０分間浸出をすることで、亜鉛の浸出率は９６．
５％を達成した。この操作に係る時間は磨鉱時間を加味
しても３２分程度であり、実施例１と同様に短時間で高
浸出率を達成することができた。By repeating the above operation twice and finally leaching for 10 minutes, the leaching rate of zinc is 96.
5% was achieved. The time required for this operation was about 32 minutes even when the grinding time was taken into account, and a high leaching rate could be achieved in a short time as in Example 1.
【００１７】[0017]
【発明の効果】ボールミル等の粉砕機を使用し、亜鉛精
鉱の粒子表面に生成する単体硫黄等の副生物を剥離、分
離しながら浸出を行うことにより、従来６〜１０時間程
度必要であった浸出時間をその１０分の１以下である３
０分程度まで短縮することが可能になった。浸出により
消費される三価の鉄イオンは、配管内に酸素を導入しな
がら再生することが可能であり、オートクレーブ等の圧
力容器が不要となった。従って、従来法に比べて浸出時
間を大幅に短縮することができ、また酸化に必要な設備
が不要となるため、これらの方法を組み合わせて利用す
ることによって建設費、操業コストの大幅な削減が可能
となった。また、本発明の方法は、既存の設備への組み
込みが可能であり、小規模の建設によって亜鉛生産量の
増産を行うことが可能になった等の効果を奏するもので
ある。According to the present invention, leaching is carried out by using a pulverizer such as a ball mill while exfoliating and separating by-products such as elemental sulfur generated on the surface of the zinc concentrate particles, which has conventionally required about 6 to 10 hours. Leaching time less than 1/10 of that
The time can be reduced to about 0 minutes. Trivalent iron ions consumed by leaching can be regenerated while introducing oxygen into the piping, and a pressure vessel such as an autoclave is not required. Therefore, the leaching time can be greatly reduced as compared with the conventional method, and the equipment required for oxidation is not required. By using these methods in combination, the construction cost and operation cost can be significantly reduced. It has become possible. Further, the method of the present invention has effects such as being able to be incorporated into existing equipment and increasing the production of zinc by small-scale construction.
───────────────────────────────────────────────────── フロントページの続き (72)発明者 鳴海 明 東京都千代田区丸の内一丁目８番２号 同 和鉱業株式会社内 Ｆターム(参考） 4K001 AA30 BA06 DB03 DB25 ────────────────────────────────────────────────── ─── Continued on the front page (72) Inventor Akira Narumi 1-8-2 Marunouchi, Chiyoda-ku, Tokyo Dowa Mining Co., Ltd. F-term (reference) 4K001 AA30 BA06 DB03 DB25 DB25
Claims (9)
水溶液中において亜鉛精鉱を磨鉱することによって該亜
鉛精鉱中の亜鉛を浸出することを特徴とする亜鉛精鉱の
浸出法。1. A method for leaching zinc concentrate, comprising leaching zinc in the zinc concentrate by grinding the zinc concentrate in an aqueous solution containing free sulfuric acid and trivalent iron ions. .
載の浸出法。2. The leaching method according to claim 1, wherein the grinding is performed under atmospheric pressure.
面に生成される副生物を前記磨鉱により剥離して該浸出
時間を短縮する、請求項１または２記載の浸出法。3. The leaching method according to claim 1, wherein by-products generated on the surface of the zinc concentrate particles along with the leaching are separated by the grinding to shorten the leaching time.
元されて生じた二価の鉄イオンを含有する浸出後液中に
酸素を供給することによって該二価の鉄イオンを三価の
鉄イオンに酸化した液を前記磨鉱の工程に繰り返す、請
求項１〜３のいずれかに記載の浸出法。4. The method according to claim 1, wherein oxygen is supplied to the leached solution containing divalent iron ions generated by reduction of said trivalent iron ions by said leaching, thereby converting said divalent iron ions to trivalent iron ions. The leaching method according to any one of claims 1 to 3, wherein the liquid oxidized into ions is repeated in the grinding step.
る前記水溶液または前記浸出後液の送液配管内に酸素を
供給して該管内を加圧状態にする、請求項１〜４のいず
れかに記載の浸出法。5. The method according to claim 1, wherein oxygen is supplied to a pipe for feeding the aqueous solution or the liquid after leaching, which is supplied or repeated to the grinding process, so that the pipe is pressurized. The leaching method described in 1.
時点で４０ｇ／L以上としてジャロサイトの生成を抑制
する、請求項１〜５のいずれかに記載の浸出法。6. The leaching method according to claim 1, wherein the concentration of free sulfuric acid in the aqueous solution is set to 40 g / L or more at the end of leaching to suppress the generation of jarosite.
〜６０ｇ／Lの範囲とする、請求項１〜６のいずれかに
記載の浸出法。7. The concentration of trivalent iron ions in the aqueous solution is 5
The leaching method according to any one of claims 1 to 6, wherein the leaching method is in a range of 60 g / L.
の金属元素を含有する前記亜鉛精鉱中の該金属元素が濃
縮された前記浸出後の残渣を浮選し該金属元素を回収す
る、請求項１〜７のいずれかに記載の浸出法。8. The leaching residue enriched with the metal element in the zinc concentrate containing at least one metal element of lead, gold and silver is flotated to recover the metal element. The leaching method according to any one of claims 1 to 7.
融点以上の温度に加熱して該浮鉱中の単体硫黄を揮発回
収する、請求項８記載の浸出法。9. The leaching method according to claim 8, wherein the flotation obtained by the flotation is heated to a temperature equal to or higher than the melting point of sulfur to volatilize and recover elemental sulfur in the flotation.
Priority Applications (8)
| Application Number | Priority Date | Filing Date | Title |
|---|---|---|---|
| JP2001091738A JP4765114B2 (en) | 2001-03-28 | 2001-03-28 | Zinc concentrate leaching method |
| US10/058,438 US6835230B2 (en) | 2001-03-28 | 2002-01-28 | Method for leaching zinc concentrate |
| DE60216346T DE60216346D1 (en) | 2001-03-28 | 2002-01-29 | Process and apparatus for leaching zinc concentrates |
| ES02002172T ES2275762T3 (en) | 2001-03-28 | 2002-01-29 | PROCEDURE AND DEVICE FOR LINIVIATION OF CONCENTRATES OF ZINC. |
| AT02002172T ATE346961T1 (en) | 2001-03-28 | 2002-01-29 | METHOD AND DEVICE FOR LEACHING ZINC CONCENTRATES |
| EP02002172A EP1245686B1 (en) | 2001-03-28 | 2002-01-29 | Method and apparatus for leaching zinc concentrates |
| KR1020020012415A KR100729192B1 (en) | 2001-03-28 | 2002-03-08 | Method and apparatus for leaching zinc concentrates |
| US10/681,455 US20040065987A1 (en) | 2001-03-28 | 2003-10-07 | Apparatus for leaching zinc concentrates |
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| Application Number | Priority Date | Filing Date | Title |
|---|---|---|---|
| JP2001091738A JP4765114B2 (en) | 2001-03-28 | 2001-03-28 | Zinc concentrate leaching method |
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| JP4765114B2 JP4765114B2 (en) | 2011-09-07 |
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Cited By (3)
| Publication number | Priority date | Publication date | Assignee | Title |
|---|---|---|---|---|
| JP2003082420A (en) * | 2001-09-13 | 2003-03-19 | Dowa Mining Co Ltd | Method and apparatus for leaching zinc concentrate |
| WO2007077290A1 (en) * | 2006-01-04 | 2007-07-12 | Outotec Oyj. | Method for improving sulphidic concentrate leaching |
| CN113462898A (en) * | 2021-06-19 | 2021-10-01 | 西部矿业股份有限公司 | Novel oxygen pressure leaching zinc smelting purification impurity removal process |
Citations (1)
| Publication number | Priority date | Publication date | Assignee | Title |
|---|---|---|---|---|
| JPH0790404A (en) * | 1990-02-16 | 1995-04-04 | Outokumpu Oy | Hydrometallurgical method for processing raw material containing zinc sulfide |
-
2001
- 2001-03-28 JP JP2001091738A patent/JP4765114B2/en not_active Expired - Lifetime
Patent Citations (1)
| Publication number | Priority date | Publication date | Assignee | Title |
|---|---|---|---|---|
| JPH0790404A (en) * | 1990-02-16 | 1995-04-04 | Outokumpu Oy | Hydrometallurgical method for processing raw material containing zinc sulfide |
Cited By (4)
| Publication number | Priority date | Publication date | Assignee | Title |
|---|---|---|---|---|
| JP2003082420A (en) * | 2001-09-13 | 2003-03-19 | Dowa Mining Co Ltd | Method and apparatus for leaching zinc concentrate |
| WO2007077290A1 (en) * | 2006-01-04 | 2007-07-12 | Outotec Oyj. | Method for improving sulphidic concentrate leaching |
| CN113462898A (en) * | 2021-06-19 | 2021-10-01 | 西部矿业股份有限公司 | Novel oxygen pressure leaching zinc smelting purification impurity removal process |
| CN113462898B (en) * | 2021-06-19 | 2022-09-06 | 西部矿业股份有限公司 | Oxygen pressure leaching zinc smelting purification impurity removal process |
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